Beneficiation process of high-silica collophanite

Abstract

The invention discloses a beneficiation process of high-silica collophanite. The beneficiation process comprises the following steps: S1, crushing; s2, screening the crushed ores to obtain ores with different size fractions; s3-S9, carrying out different silicon reduction processes such as photoelectric separation, dense medium and gravity separation on the ores with different particle sizes according to proper particle sizes; after respective tailings of different size fractions are scavenged, obtained scavenged concentrates are returned, and the obtained tailings serve as pre-enriched tailings; s10, the fine-fraction ore and the pre-enriched concentrate are combined and ground, and to-be-separated ore pulp is obtained; s7, water is added into the to-be-separated ore pulp, flotation ore pulp is obtained, and final phosphate concentrate and final tailings are obtained through one-rough-one-fine-one-scavenging flotation operation; through step-by-step silicon reduction, the beneficiation efficiency is high, the amount of floated ores is small, the energy consumption is low, the cost is low, and the beneficiation method is environmentally friendly.

Description

Technical Field

[0001] The invention relates to the technical field of collophanite beneficiation, in particular to a high-silica collophanite beneficiation process. Background Art

[0003] The current mineral processing technology is mainly based on direct and reverse flotation technology or double reverse flotation technology, which has a series of problems such as long process, complex medium, high reagent consumption, and difficult tailings water treatment, resulting in high mineral processing costs, low mineral processing efficiency, and high environmental pollution.

[0004] In summary, the existing technology of flotation mixed type collophanite beneficiation has the problems of high cost and large amount of reagents.

[0005] Therefore, we urgently need a high-silica collophanite beneficiation process that is simple to operate, low in cost and environmentally friendly while improving the beneficiation efficiency. Summary of the invention

[0006] The purpose of the present invention is to overcome the shortcomings of the prior art and provide a high-silica collophanite beneficiation process, including photoelectric separation, heavy medium separation and single reverse flotation process. Through photoelectric separation, heavy medium separation to reduce silicon and flotation to remove impurities, the effects of high beneficiation efficiency, small amount of ore to be floated, low energy consumption, low flotation agent cost and environmental friendliness are achieved.

[0007] The objective of the present invention is achieved through the following technical solutions:

[0008] The present invention provides a high-silica collophanite beneficiation process for step-by-step silicon reduction and impurity removal, comprising the following steps:

[0009] S1. crushing the raw ore to obtain crushed ore with a particle size of ≤30mm;

[0010] S2. The crushed ore is screened to obtain ore with a particle size of -30mm+15mm and ore with a particle size of -30mm-15mm;

[0011] S3. Use photoelectric separation technology for ores with a particle size of -30mm+15mm to obtain photoelectric concentrate and photoelectric tailings;

[0012] S4. The photoelectric tailings are subjected to photoelectric scavenging to obtain photoelectric scavenging concentrate and photoelectric scavenging tailings, respectively. The photoelectric scavenging concentrate is returned to S3 operation, and the photoelectric scavenging tailings with a P2O5 grade of <10% are selected as the first tailings;

[0013] S5. The photoelectric concentrate obtained in S3 is further crushed and merged into the -30mm-15mm ore obtained in S2, and the merged ore is screened and classified to obtain -15mm+0.5mm ore and -15mm-0.5mm ore; wherein the -15mm+0.5mm ore is sorted to obtain a heavy medium concentrate and a heavy medium tailings;

[0014] S6. The heavy medium tailings obtained in S5 are subjected to heavy medium scavenging to obtain heavy medium scavenging concentrate and heavy medium scavenging tailings, the heavy medium scavenging concentrate is returned to S5, and the heavy medium scavenging tailings with a P2O5 grade of <10% are selected as the second tailings;

[0015] S7. The heavy medium concentrate obtained in the S5 operation is further crushed and combined with the -15mm-0.5mm ore obtained in S5. The combined ore is screened and classified to obtain -0.5mm+0.038mm ore and -0.5mm-0.038mm ore; wherein the -0.5mm+0.038mm ore is re-selected to obtain a re-selection concentrate and a re-selection tailings;

[0016] S8. The gravity separation tailings obtained in S7 are scavenged to obtain gravity separation scavenging concentrate and gravity separation scavenging tailings, the gravity separation scavenging concentrate is returned to S7 operation, and the gravity separation scavenging tailings with a P2O5 grade of <10% are selected as the third tailings;

[0017] S9. Combine the -0.5mm-0.038mm particle size ore and the re-selection scavenging concentrate obtained in S8 to obtain a pre-enriched concentrate; combine the first tailings obtained in S4, the second tailings obtained in S6 and the third tailings obtained in S8 to obtain a pre-enriched tailings;

[0018] S10. Grinding the pre-enriched concentrate to obtain flotation pulp, and subjecting the flotation pulp to a coarse-fine-sweep operation to obtain a final phosphate concentrate and a final tailings.

[0019] Through the above technical scheme, the photoelectric separation, heavy medium and gravity separation combined separation process is adopted to realize the desiliconization and tailings discarding of high-siliceous collophosphate ore, avoiding the influence of siliceous gangue on subsequent flotation operations, reducing the crushing and grinding processing volume, reducing the flotation processing volume, improving the flotation efficiency, and saving production costs.

[0020] Furthermore, in S10, in the flotation slurry, the weight of minerals with a particle size of ≤0.074 mm accounts for 75% to 90% of the total weight. At this time, the mineral particles are most likely to combine with the reagent molecules to form effective mineralized foam to complete the separation.

[0021] Further, in S10, the one-rough-one-fine-one-sweep flotation operation comprises the following steps:

[0022] The one rough one fine one sweeping operation comprises the following steps:

[0023] A1. Add 1800-2200 g / t of the original ore inhibitor and 800-1000 g / t of the original ore collector to the flotation slurry, and obtain the rougher concentrate and rougher tailings by stirring and aerating;

[0024] A2. Adding 400-500 g / t of the raw ore as an inhibitor and 200-300 g / t of the raw ore as a collector to the rougher concentrate, stirring and aerating to obtain the final phosphate concentrate and the selected middlings;

[0025] A3. Add 900-1100 g / t of inhibitor and 400-500 g / t of collector to the rougher tailings, and obtain scavenger concentrate and final tailings by stirring and aerating.

[0026] Through the above technical solution, the intermediate products in the one-rough-one-fine-one-sweep flotation operation are further separated, thereby achieving the effect of improving the yield of the final phosphate concentrate.

[0027] Through the above technical scheme, since the amphoteric adjuvant is a collector molecule containing amino and carboxyl groups, the amino group is a cation, and under suitable pH conditions, it can form electrostatic adsorption with the surface of the mineral with negative charge. The carboxyl group is a functional group of the anionic collector, which can form chemical adsorption with the metal ions with positive charge on the surface of the mineral. Through the adsorption of the hydrophilic group anion and cation on the mineral surface, the collector molecule forms an effective hydrophobic layer on the mineral surface, thereby enhancing the collector's ability to capture dolomite and siliceous minerals in the slurry, achieving improved selectivity and capture performance, and ensuring that apatite minerals and dolomite and siliceous gangue minerals are effectively separated.

[0028] Furthermore, the inhibitor includes 2 to 3 parts of sulfuric acid, 1 to 2 parts of phosphoric acid, 1 to 2 parts of hydroxyethylene diphosphoric acid and 2 to 3 parts of sodium hexametaphosphate in parts by weight.

[0029] Furthermore, the preparation method of the collector comprises:

[0030] After mixing the amphoteric adjuvant, octadecyl phosphate, lauric acid and vegetable oil, a mixed solution is obtained; and NaOH is added to the mixed solution to perform a first heating reaction to obtain the collector;

[0031] Wherein, by weight, the amphoteric adjuvant is 4-5 parts, the octadecyl phosphate is 1 part, the lauric acid is 1 part, and the vegetable oil is 3-4 parts.

[0032] Furthermore, the weight ratio of the NaOH to the mixed solution is (0.1-0.2):1.

[0033] Furthermore, the preparation method of the amphoteric adjuvant comprises the following steps: adding methyl linoleate to dodecyl primary amine, stirring and then performing a second heating reaction to obtain the amphoteric adjuvant;

[0034] Wherein, the weight ratio of the dodecyl primary amine to the methyl linoleate is 1:(3-4);

[0035] The temperature of the second heating reaction is 100 to 120° C., and the time of the second heating reaction is 40 to 50 minutes.

[0036] Furthermore, the vegetable oil is a mixture of cottonseed oil and coconut oil.

[0037] Furthermore, the weight ratio of the cottonseed oil to the coconut oil is (1.5-2.5):1.

[0038] Furthermore, the temperature of the first heating reaction is 60-80° C., and the time of the first heating reaction is 3-5 hours.

[0039] The beneficial effects of the present invention are:

[0040] 1. The high-siliceous collophanite beneficiation process of the present invention combines multiple processes to remove impurities and reduce silicon, including pre-selection desiliconization by photoelectric separation and heavy medium separation, mixed desiliconization by amphoteric adjuvants, and different desiliconization and impurity removal processes are used for siliceous minerals in different embedded states, thereby achieving a low-energy and high-efficiency sorting effect.

[0041] 2. The high-silica collophosphate beneficiation process of the present invention removes impurities from gangue minerals under coarse particle size conditions, thereby achieving the effects of significantly reducing grinding energy consumption and flotation agent consumption, effectively reducing production energy consumption and saving production costs.

[0042] 3. The high-siliceous collophosphate beneficiation process of the present invention increases the flotation of fine-grained siliceous minerals in the flotation slurry by adding an amphoteric adjuvant to the collector, so that the collector molecules can be more effectively adsorbed on the surface of calcium-magnesium and siliceous minerals, thereby achieving the effect of effectively improving the sorting property of apatite minerals. BRIEF DESCRIPTION OF THE DRAWINGS

[0043] Figure 1 The present invention is a flow chart of a high-silica collophanite beneficiation process. DETAILED DESCRIPTION

[0044] The technical solution of the present invention is further described in detail below in conjunction with the accompanying drawings, but the protection scope of the present invention is not limited to the following.

[0045] Example 1

[0046] Prepare the amphoteric adjuvant as follows:

[0047] S11. Add 3 parts of methyl linoleate dropwise into 1 part of dodecyl primary amine while stirring, and heat to 100° C. and react for 40 minutes to obtain the amphoteric adjuvant.

[0048] Example 2

[0049] Prepare the amphoteric adjuvant as follows:

[0050] S11. 3.5 parts of methyl linoleate are added dropwise to 1 part of dodecyl primary amine while stirring, and the mixture is heated to 110° C. and reacted for 45 minutes to obtain the amphoteric adjuvant.

[0051] Example 3

[0052] Prepare the amphoteric adjuvant as follows:

[0053] S11. 4 parts of methyl linoleate are added dropwise to 1 part of dodecyl primary amine, while stirring, and heated to 120° C., and reacted for 50 minutes to obtain the amphoteric adjuvant.

[0054] Example 4

[0055] Prepare vegetable oils as follows:

[0056] S21. Mix 3 parts of cottonseed oil and 2 parts of coconut oil to obtain the vegetable oil.

[0057] Example 5

[0058] Prepare vegetable oils as follows:

[0059] S21. Mix 3 parts of cottonseed oil and 2 parts of coconut oil to obtain the vegetable oil.

[0060] Example 6

[0061] Prepare vegetable oils as follows:

[0062] S21. Mix 5 parts of cottonseed oil and 2 parts of coconut oil to obtain the vegetable oil.

[0063] Example 7

[0064] Prepare the collector as follows:

[0065] S31. After mixing 4 parts of the amphoteric adjuvant prepared in Example 1, 1 part of octadecyl phosphate, 1 part of lauric acid and 4 parts of the vegetable oil prepared in Example 4, a mixed solution was obtained;

[0066] S22. Add 1 part of NaOH to 10 parts of the mixed solution, heat to 60°C and react for 5 hours to obtain the collector.

[0067] Example 8

[0068] Prepare the collector as follows:

[0069] S31. After mixing 5 parts of the amphoteric adjuvant prepared in Example 1, 1 part of octadecyl phosphate, 1 part of lauric acid and 3 parts of the vegetable oil prepared in Example 5, a mixed solution was obtained;

[0070] S32. Add 1.5 parts of NaOH to 10 parts of the mixed solution, heat to 70°C and react for 4 hours to obtain the collector.

[0071] Example 9

[0072] Prepare the collector as follows:

[0073] S31. After mixing 4 parts of the amphoteric adjuvant prepared in Example 1, 1 part of octadecyl phosphate, 1 part of lauric acid and 4 parts of the vegetable oil prepared in Example 6, a mixed solution was obtained;

[0074] S32. Add 2 parts of NaOH to 10 parts of the mixed solution, heat to 80°C and react for 3 hours to obtain the collector.

[0075] Example 10

[0076] A high-siliceous collophosphate ore, taken from a concentrator in Leibo area, with a P2O5 grade of 20%. Its concentrating process is as follows: Figure 1 As shown, the following steps are included:

[0077] S41. Crushing the raw ore to obtain crushed ore with a particle size of ≤30 mm;

[0078] S42. The crushed ore is screened to obtain ores of -30mm+15mm and -30mm-15mm particle sizes, respectively;

[0079] S43. Photoelectric separation process is used for -30mm+15mm particle size ore to obtain photoelectric concentrate and photoelectric tailings;

[0080] S44. The photoelectric tailings obtained are subjected to photoelectric scavenging separation to obtain photoelectric scavenging concentrate and scavenging tailings respectively. The scavenging concentrate is returned to S43 operation, and the photoelectric scavenging tailings with P2O5 grade <10% are used as the first tailings.

[0081] S45. The photoelectric concentrate obtained in S43 is further crushed and merged into the -15mm particle size obtained in S42. The obtained mixed -15mm particle size is screened and classified to obtain -15mm+0.5mm and -15mm-0.5mm particle sizes; among which, the -15mm+0.5mm particle size is separated by a heavy medium cyclone to obtain a heavy medium concentrate and a heavy medium tailings.

[0082] The heavy medium tailings obtained from S46.S45 are subjected to heavy medium scavenging to obtain heavy medium scavenging concentrate and scavenging tailings. The scavenging concentrate is returned to S45 operation as a roughing feed for the heavy medium cyclone, and the heavy medium scavenging tailings with a P2O5 grade of <10% is used as the second tailings.

[0083] The heavy medium concentrate obtained from the S47.S45 operation is further crushed and merged into the -0.5mm particle size obtained from S45. The obtained mixed -0.5mm particle size is screened and graded to obtain -0.5mm+0.038mm and -0.038mm particle sizes; among them, the -0.5mm+0.038mm particle size is re-selected by a spiral chute to obtain re-selection concentrate and re-selection tailings.

[0084] The gravity separation tailings obtained from S48 and S47 are scavenged by spiral chute to obtain gravity separation scavenging concentrate and scavenging tailings. The scavenging concentrate is returned to S47 operation as the spiral chute roughing feed, and the gravity separation scavenging tailings with P2O5 grade <10% are used as the third tailings.

[0085] S49. Combine the -0.038mm particle size and the gravity separation concentrate obtained by S48 to obtain a pre-enriched concentrate; combine the photoelectric separation tailings obtained by S44, the heavy medium tailings obtained by S46 and the gravity separation tailings obtained by S48 to obtain a pre-enriched tailings.

[0086] S410. Grind the pre-enriched concentrate to obtain flotation pulp, and adopt a coarse-fine-sweep flotation operation to obtain the final phosphate concentrate and the final tailings. Specifically, the following steps are included:

[0087] A1. Add 2000 g / t of the original ore inhibitor and 800 g / t of the original ore collector prepared in Example 7 to the flotation slurry (the mineral with a particle size of ≤0.074 mm accounts for 75% of the total weight), and stir and aerate to obtain a rougher concentrate and a rougher tailings;

[0088] A2. Adding 500 g / t of the raw ore inhibitor and 200 g / t of the raw ore collector prepared in Example 7 to the rougher concentrate, stirring and aerating to obtain the final phosphate concentrate and selected ore;

[0089] A3. Adding 1000 g / t of the inhibitor of the raw ore and 400 g / t of the raw ore of the collector prepared in Example 7 to the rougher tailings, stirring and aerating to obtain a scavenged concentrate and a final tailings;

[0090] The selected middlings and scavenged concentrates can be returned to step A1 respectively, and steps A1 to A3 are repeated;

[0091] In parts by weight, the inhibitor includes 3 parts of sulfuric acid, 2 parts of phosphoric acid, 2 parts of hydroxyethylene diphosphoric acid and 3 parts of sodium hexametaphosphate.

[0092] Embodiment 11

[0093] A high-siliceous collophosphate ore, obtained from a concentrator in Mabian area, with a P2O5 grade of 19%. Its concentrating process is as follows: Figure 1 As shown, the following steps are included:

[0094] S41. Crushing the raw ore to obtain crushed ore with a particle size of ≤30 mm;

[0095] S42. The crushed ore is screened to obtain ores of different particle sizes such as -30mm+15mm and -30mm-15mm;

[0096] S43. Photoelectric separation process is used for -30mm+15mm particle size ore to obtain photoelectric concentrate and photoelectric tailings;

[0097] S44. The photoelectric tailings obtained are subjected to photoelectric scavenging separation to obtain photoelectric scavenging concentrate and scavenging tailings respectively. The scavenging concentrate is returned to S43 operation, and the photoelectric scavenging tailings with P2O5 grade <10% are used as the first tailings.

[0098] S45. The photoelectric concentrate obtained in S43 is further crushed and merged into the -15mm particle size obtained in S42. The obtained mixed -15mm particle size is screened and classified to obtain -15mm+0.5mm and -0.5mm particle sizes; among which, the -15mm+0.5mm particle size is separated by a heavy medium cyclone to obtain a heavy medium concentrate and a heavy medium tailings.

[0099] The heavy medium tailings obtained from S46.S45 are subjected to heavy medium scavenging to obtain heavy medium scavenging concentrate and scavenging tailings. The scavenging concentrate is returned to S45 operation as a roughing feed for the heavy medium cyclone, and the heavy medium scavenging tailings with a P2O5 grade of <10% is used as the second tailings.

[0100] The heavy medium concentrate obtained from the S47.S45 operation is further crushed and merged into the -0.5mm particle size obtained from S45. The obtained mixed -0.5mm particle size is screened and graded to obtain -0.5mm+0.038mm and -0.038mm particle sizes; among them, the -0.5mm+0.038mm particle size is re-selected by a spiral chute to obtain re-selection concentrate and re-selection tailings.

[0101] The gravity separation tailings obtained from S48 and S47 are scavenged by spiral chute to obtain gravity separation scavenging concentrate and scavenging tailings. The scavenging concentrate is returned to S47 operation as the spiral chute roughing feed, and the gravity separation scavenging tailings with P2O5 grade <10% are used as the third tailings.

[0102] S49. Combine the -0.038mm particle size and the gravity separation concentrate obtained by S48 to obtain a pre-enriched concentrate; combine the photoelectric separation tailings obtained by S44, the heavy medium tailings obtained by S46 and the gravity separation tailings obtained by S48 to obtain a pre-enriched tailings.

[0103] S410. Grind the pre-enriched concentrate to obtain flotation pulp, and adopt a coarse-fine-sweep flotation operation to obtain the final phosphate concentrate and the final tailings, which specifically includes the following steps:

[0104] A1. Add 1800 g / t of the original ore inhibitor and 1000 g / t of the original ore collector prepared in Example 8 to the flotation slurry (the mineral with a particle size of ≤0.074 mm accounts for 82% of the total weight), and stir and aerate to obtain a rougher concentrate and a rougher tailings;

[0105] A2. Adding 400 g / t of the raw ore inhibitor and 300 g / t of the raw ore collector prepared in Example 8 to the rougher concentrate, stirring and aerating to obtain the final phosphate concentrate and selected ore;

[0106] A3. Adding 900 g / t of the inhibitor of the raw ore and 500 g / t of the raw ore of Example 8 to the rougher tailings, stirring and aerating to obtain a scavenged concentrate and a final tailings;

[0107] The selected middlings and scavenged concentrates can be returned to step A1 respectively, and steps A1 to A3 are repeated;

[0108] In parts by weight, the inhibitor includes 2 parts of sulfuric acid, 1 part of phosphoric acid, 1 part of hydroxyethylene diphosphoric acid and 3 parts of sodium hexametaphosphate.

[0109] Example 12

[0110] A high-siliceous collophosphate ore, obtained from a concentrator in Jinyang, with a P2O5 grade of 18%. Figure 1 As shown, the following steps are included:

[0111] S41. Crushing the raw ore to obtain crushed ore with a particle size of ≤30 mm;

[0112] S42. The crushed ore is screened to obtain ores of different particle sizes such as -30mm+15mm and -15mm;

[0113] S43. Photoelectric separation process is used for -30mm+15mm particle size ore to obtain photoelectric concentrate and photoelectric tailings;

[0114] S44. The photoelectric tailings obtained are subjected to photoelectric scavenging separation to obtain photoelectric scavenging concentrate and scavenging tailings respectively. The scavenging concentrate is returned to S43 operation, and the photoelectric scavenging tailings with P2O5 grade <10% are used as the first tailings.

[0115] S45. The photoelectric concentrate obtained in S43 is further crushed and merged into the -15mm particle size obtained in S42. The obtained mixed -15mm particle size is screened and classified to obtain -15mm+0.5mm and -0.5mm particle sizes; among which, the -15mm+0.5mm particle size is separated by a heavy medium cyclone to obtain a heavy medium concentrate and a heavy medium tailings.

[0116] The heavy medium tailings obtained from S46.S45 are subjected to heavy medium scavenging to obtain heavy medium scavenging concentrate and scavenging tailings. The scavenging concentrate is returned to S45 operation as a roughing feed for the heavy medium cyclone, and the heavy medium scavenging tailings with a P2O5 grade of <10% is used as the second tailings.

[0117] The heavy medium concentrate obtained from the S47.S45 operation is further crushed and merged into the -0.5mm particle size obtained from S45. The obtained mixed -0.5mm particle size is screened and graded to obtain -0.5mm+0.038mm and -0.038mm particle sizes; among them, the -0.5mm+0.038mm particle size is re-selected by a spiral chute to obtain re-selection concentrate and re-selection tailings.

[0118] The gravity separation tailings obtained from S48 and S47 are scavenged by spiral chute to obtain gravity separation scavenging concentrate and scavenging tailings. The scavenging concentrate is returned to S47 operation as the spiral chute roughing feed, and the gravity separation scavenging tailings with P2O5 grade <10% are used as the third tailings.

[0119] S49. Combine the -0.038mm particle size and the gravity separation concentrate obtained by S48 to obtain a pre-enriched concentrate; combine the photoelectric separation tailings obtained by S44, the heavy medium tailings obtained by S46 and the gravity separation tailings obtained by S48 to obtain a pre-enriched tailings.

[0120] S410. Grind the pre-enriched concentrate to obtain flotation pulp, and adopt a coarse-fine-sweep flotation operation to obtain the final phosphate concentrate and the final tailings, which specifically includes the following steps:

[0121] A1. Add 2200 g / t of the original ore inhibitor and 1000 g / t of the original ore collector prepared in Example 9 to the flotation slurry (the mineral with a particle size of ≤0.074 mm accounts for 90% of the total weight), and stir and aerate to obtain a rougher concentrate and a rougher tailings;

[0122] A2. Adding 500 g / t of the raw ore inhibitor and 300 g / t of the raw ore collector prepared in Example 9 to the rougher concentrate, stirring and aerating to obtain the final phosphate concentrate and selected ore;

[0123] A3. Adding 1100 g / t of the inhibitor of the raw ore and 500 g / t of the raw ore of Example 9 to the rougher tailings, stirring and aerating to obtain a scavenged concentrate and a final tailings;

[0124] The selected middlings and scavenged concentrates can be returned to step A1 respectively, and steps A1 to A3 are repeated;

[0125] In parts by weight, the inhibitor includes 2 parts of sulfuric acid, 2 parts of phosphoric acid, 2 parts of hydroxyethylene diphosphoric acid and 2 parts of sodium hexametaphosphate.

[0126] Comparative Example 1

[0127] The various indicators of separating high-silica collophanite in Example 10 of the present invention were compared with those in Control Example 1, wherein Control Example 1 used the medium- and low-grade mixed collophanite in Example 10, and its mineral processing technology was: the technical solution described in Chinese patent document CN201510991054.3.

[0128] Comparative Example 2

[0129] The various indicators of separating high-silica collophanite in Example 10 of the present invention were compared with those in Control Example 2, wherein Control Example 2 used the collophanite in Example 10, and its beneficiation process was as follows: removing steps S41 to S410 in Example 10, directly grinding the original ore to obtain the ore pulp to be separated, and other conditions such as the amount of reagents used, the grinding fineness and the subsequent process flow adopted were the same as those in Example 10 of the present invention.

[0130] Comparative Example 3

[0131] The various indicators of separating high-silica collophanite in Example 10 of the present invention are compared with those in Control Example 3, wherein Control Example 3 uses the high-silica collophanite in Example 10, and its beneficiation process is: the collector is replaced with a fatty acid salt mainly composed of sodium oleate in the prior art, and other conditions such as the dosage of the reagent, the grinding fineness and the adopted process flow are the same as those in Example 10 of the present invention. (This control example is compared with the case where the collector is replaced with the prior art at the same time, to prove that the beneficiation process of the present invention has a better effect).

[0132] The experimental results of Examples 10 to 12 and Comparative Examples 1 to 3 are compared as shown in Table 1 below (wherein, the raw ore grade in Comparative Examples 1 to 3 is slightly different from that in Example 10 because of human error in the experimental operation, which is a normal phenomenon):

[0133] Table 1 Comparison results of Examples 10 to 12 and Comparative Examples 1 to 3

[0134]

[0135]

[0136] As can be seen from the above table, compared with Example 10, the yield, grade and recovery rate of the final phosphate concentrate obtained in Reference Example 1 are significantly reduced, while the yield, grade and recovery rate of the obtained tailings are significantly increased; the yield of the final phosphate concentrate obtained in Reference Example 2 is slightly reduced, the grade and recovery rate are significantly reduced, while the yield of the obtained tailings is slightly increased, and the grade and recovery rate are significantly increased; the yield of the final phosphate concentrate obtained in Reference Example 3 is increased, and the grade is significantly reduced, while the yield of the obtained tailings is decreased, and the grade is significantly increased. It can be seen that Example 10 of the present invention can more fully separate the phosphate concentrate from the raw ore compared with the methods of Reference Examples 1 to 3; at the same time, although the yield of Reference Example 3 is increased, its grade is significantly reduced, which means that other impurities float together, proving that the collector performance of Reference Example 3 is poor compared with Example 10.

[0137] Therefore, the high-silica collophanite beneficiation process of the present invention achieves the effects of high beneficiation efficiency, small amount of ore to be floated, low energy consumption, low flotation reagent cost and environmental friendliness.

[0138] The above is only a preferred embodiment of the present invention. It should be understood that the present invention is not limited to the form disclosed herein, and should not be regarded as excluding other embodiments, but can be used in various other combinations, modifications and environments, and can be modified within the scope of the concept described herein through the above teachings or the technology or knowledge of the relevant field. The changes and modifications made by those skilled in the art shall not deviate from the spirit and scope of the present invention, and shall be within the scope of protection of the claims attached to the present invention.

Claims

1. A high-silica collophanite beneficiation process, characterized in that: The following steps are involved: S1. crushing the raw ore to obtain crushed ore with a particle size of ≤30mm; S2. The crushed ore is screened to obtain ore with a particle size of -30mm+15mm and ore with a particle size of -30mm-15mm; S3. Use photoelectric separation technology for ores with a particle size of -30mm+15mm to obtain photoelectric concentrate and photoelectric tailings; S4. The photoelectric tailings are subjected to photoelectric scavenging to obtain photoelectric scavenging concentrate and photoelectric scavenging tailings, respectively. The photoelectric scavenging concentrate is returned to S3 operation, and the photoelectric scavenging tailings with a P2O5 grade of <10% are selected as the first tailings; S5. The photoelectric concentrate obtained in S3 is further crushed and merged into the -30mm-15mm ore obtained in S2, and the merged ore is screened and classified to obtain -15mm+0.5mm ore and -15mm-0.5mm ore; wherein the -15mm+0.5mm ore is sorted to obtain a heavy medium concentrate and a heavy medium tailings; S6. The heavy medium tailings obtained in S5 are subjected to heavy medium scavenging to obtain heavy medium scavenging concentrate and heavy medium scavenging tailings, the heavy medium scavenging concentrate is returned to S5, and the heavy medium scavenging tailings with a P2O5 grade of <10% are selected as the second tailings; S7. The heavy medium concentrate obtained in the S5 operation is further crushed and combined with the -15mm-0.5mm ore obtained in S5. The combined ore is screened and classified to obtain -0.5mm+0.038mm ore and -0.5mm-0.038mm ore; wherein the -0.5mm+0.038mm ore is re-selected to obtain a re-selection concentrate and a re-selection tailings; S8. The gravity separation tailings obtained in S7 are scavenged to obtain gravity separation scavenging concentrate and gravity separation scavenging tailings, the gravity separation scavenging concentrate is returned to S7 operation, and the gravity separation scavenging tailings with a P2O5 grade of <10% are selected as the third tailings; S9. Combine the -0.5mm-0.038mm particle size ore and the re-selection scavenging concentrate obtained in S8 to obtain a pre-enriched concentrate; combine the first tailings obtained in S4, the second tailings obtained in S6 and the third tailings obtained in S8 to obtain a pre-enriched tailings; S10. Grinding the pre-enriched concentrate to obtain flotation pulp, and subjecting the flotation pulp to a coarse-fine-sweep operation to obtain a final phosphate concentrate and a final tailings.

2. The beneficiation process of a high-silica collophanite according to claim 1, characterized in that: In S10, in the flotation slurry, the weight of minerals with a particle size of ≤0.074 mm accounts for 75% to 90% of the total weight.

3. The beneficiation process of a high-silica collophanite according to claim 1, characterized in that: In S10, the one rough one fine one sweeping operation includes the following steps: A1. Add 1800-2200 g / t of the original ore inhibitor and 800-1000 g / t of the original ore collector to the flotation slurry, and obtain the rougher concentrate and rougher tailings by stirring and aerating; A2. Adding 400-500 g / t of the raw ore as an inhibitor and 200-300 g / t of the raw ore as a collector to the rougher concentrate, stirring and aerating to obtain the final phosphate concentrate and the selected middlings; A3. Add 900-1100 g / t of inhibitor and 400-500 g / t of collector to the rougher tailings, and obtain scavenger concentrate and final tailings by stirring and aerating.

4. The beneficiation process of high-silica collophanite according to claim 3, characterized in that: In parts by weight, the inhibitor comprises 2 to 3 parts of sulfuric acid, 1 to 2 parts of phosphoric acid, 1 to 2 parts of hydroxyethylene diphosphoric acid and 2 to 3 parts of sodium hexametaphosphate.

5. The beneficiation process of high-silica collophanite according to claim 3, characterized in that: The preparation method of the collector comprises the following steps: Mixing an amphoteric adjuvant, octadecyl phosphate, lauric acid and vegetable oil to obtain a mixed solution; and Adding NaOH to the mixed solution to perform a first heating reaction to obtain the collector; Wherein, by weight, the amphoteric adjuvant is 4-5 parts, the octadecyl phosphate is 1 part, the lauric acid is 1 part, and the vegetable oil is 3-4 parts.

6. The beneficiation process of high-silica collophanite according to claim 5, characterized in that: The weight ratio of the NaOH to the mixed solution is (0.1-0.2):

1.

7. The beneficiation process of high-silica collophanite according to claim 5, characterized in that: The preparation method of the amphoteric adjuvant comprises the following steps: Adding methyl linoleate to dodecyl primary amine, stirring and then performing a second heating reaction to obtain the amphoteric adjuvant; Wherein, the weight ratio of the dodecyl primary amine to the methyl linoleate is 1:(3-4); The temperature of the second heating reaction is 100 to 120° C., and the time of the second heating reaction is 40 to 50 minutes.

8. The beneficiation process of high-silica collophanite according to claim 5, characterized in that: The vegetable oil is prepared by mixing cottonseed oil and coconut oil.

9. The beneficiation process of high-silica collophanite according to claim 8, characterized in that: The weight ratio of the cottonseed oil to the coconut oil is (1.5-2.5):

1.

10. The beneficiation process of high-silica collophanite according to claim 5, characterized in that: The temperature of the first heating reaction is 60-80° C., and the time of the first heating reaction is 3-5 hours.

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